In order to obtain zinc metal from its raw materials, mainly zinc sulphide concentrates, both the pyrometallurgical and the hydrometallurgical routes have been used, although the first of these is clearly coming into disuse due to the high operating costs and the environmental problems associated to this process. Hydrometallurgical processes mostly follow the RLE line (Roasting, Leaching, Electrowinning), although some plants, very few, avoid roasting the concentrates, either because they carry out the direct leaching of concentrate process under pressure in autoclaves, or at atmospheric pressure.
Until the mid '60s, electrolytic zinc plants used a neutral leaching stage and a weak acid leaching stage in the leaching area. This method allowed them to extract the zinc contained in oxide form in the calcine, the product resulting from roasting, while the zinc combined with the iron in the form of zinc ferrites was not leached. This process yielded zinc recovery rates of between 85 and 90%, leaving behind a residue in which the zinc ferrites were concentrated, with a zinc content of 17-20%.
In 1965 the process known as jarosite process began to be used at the industrial level, as described in documents ES 34601, ES 385575 and NO 108047. Implementing this process was an important step towards successfully increasing the recovery rate of zinc above 90% levels. In addition to neutral leaching, the process also entails two or more stages of acid leaching where solubilizing the zinc and iron contained in zinc ferrites produces zinc sulphate (ZnSO4) and ferric sulphate (Fe2(SO4)3), while allowing at the same time to separate a residue containing the lead and silver present in the calcine. Afterwards, this solution containing Fe+++ in sulphate form and having the residual acidity necessary for keeping the Fe+++ in solution, is treated with calcine in the presence of a cation such as Na+, K+ or NH4+ under certain conditions required to partially lower the acidity and facilitating the iron to precipitate as jarosite, a basic sulphate having the formula Me(SO4)2Fe3(OH)6, where Me can be one of the cations mentioned above. Later on, incorporating a jarosite acid washing stage made possible to increase the recovery rate up to 97%. This process is efficient and its operating cost is very competitive.
A variation on the jarosite process is what is known as the conversion process described in document CA 1094819. This process differs from the process described above in that both the leaching of ferrites and the precipitation of iron as jarosite take place simultaneously, although in this process it is not possible to separate the lead-silver residue, obtaining at the end a single residue containing all the iron in the form of jarosite as well as the lead, silver and silica contained in the calcine.
Another variation of the jarosite process is described in document U.S. Pat. No. 4,305,914 of Dec. 15, 1981. It is a procedure in which the solution obtained from the acid leaching stage containing the Fe+++ in solution is cooled, and after the acidity present has been later partially neutralized, the solution is reheated again to precipitate the jarosite in the presence of a cation such as Na+, K+ or NH4+, after having diluting it with a zinc sulphate solution in order to prevent the acidity generated by precipitating the iron present as jarosite to be so high that it prevents the precipitation process. This sequence eliminates the need to neutralize with calcine, obtaining a jarosite residue with low heavy metal content. Nevertheless, this process is not cost effective and therefore has never been developed at the industrial level.
Another process developed several years after the jarosite process, known as the goethite process, is described in document CA 873262. As in the case of the jarosite process, this process entails a neutral leaching stage and one or more stages of acid leaching working in counter-current, and where the ferrites are leached while at the same time is possible to separate the lead-silver residue. The solution resulting from the acid leaching is treated with zinc concentrate in order to reduce the ferric iron (Fe+++) to ferrous iron (Fe++). This is followed by a pre-neutralization stage, where part of the existing acidity is neutralized with calcine, and a subsequent iron oxidation and precipitation stage that results in goethite (FeO(OH)), in which calcine is also used to neutralize the acidity generated in the formation of goethite and oxygen is used for oxidizing Fe++ to Fe+++. This process produces a residue that is somewhat richer in iron, between 30 and 40%, compared to the percentage obtained with the jarosite process, in which the iron content of the residue obtained is usually between 28 and 32%. However, the zinc recovery rate of this process is lower than the one obtained with the jarosite process. While the usual final zinc content found in the residue resulting from the jarosite process is usually of 3-4% of zinc, the final residue resulting from the goethite process contains up to 8-10% of zinc.
A variation of the goethite process using paragoethite yields results similar to those described above.
Nowadays there are a certain number of electrolytic zinc plants combining the traditional (RLE) process with direct leaching of concentrates. It is usual in these plants to generate a final residue containing the iron (in most cases in the form of jarosite) and also the lead, silver and silica contained in the treated raw materials in addition to the elemental sulphur generated during the direct leaching process.
The main drawbacks of these processes can be summarized below:                Zinc recovery rates, though acceptable, in the best of cases does not exceed 97%, while in the majority of plants using these processes the overall recovery ranges between 94 and 96.5%.        The percentage of lead and silver recovered with the lead-silver residue does not generally exceed 60-70% of the total of these metals contained in the calcine; in many of those plants the recovery rate for these metals is frequently around 50%. The remaining content is lost together with the iron residue, thereby contaminating it.        The recovery rate for copper does not exceed 80%, since the iron residue contains appreciable quantities of this metal.        The amount of impurities accompanying the iron residue, jarosite, goethite or paragoethite (zinc and lead as already mentioned, as well as arsenic and/or copper when zinc concentrates are treated with appreciable contents of these elements) means that the residue cannot be used for any other process and has to be stored in safety ponds, becoming a major environmental liability. In the case of jarosite, environmental regulations do not allow it to be stored in the form it is generated by the zinc manufacturing process, and therefore it has to be first rendered inert by mixing it with lime and cement (jarofix process), before it can be stored in safety ponds.        Currently, certain countries have already banned the practice of storing this kind of residue (Netherlands, Japan, Australia), while another group of countries allow it to be stored in existing ponds but no longer permit the construction of new storage ponds (France, Belgium, Germany). This situation is becoming more restrictive as environmental pressure grows, demanding cleaner and more efficient technologies for the electrolytic zinc production.        
Consequently, any novel technology intended to be applied in this field would have to enable maximum metal recovery rates at a competitive cost and generate only environmentally acceptable residues that can be in turn used favorably in other industrial processes, eliminating the need for permanent storage—a solution, as noted above, that is no longer permitted in some countries, and which will be, presumably, also banned, in other countries in a not so distant future—. In this regard, during the last 30 years intensive research work has been conducted in the field of zinc production searching for a manageable and economically competitive process which has a high metal recovery rate, although to date, no satisfactory solution has been found. One of the many examples regarding these works that can be cited is described in document U.S. Pat. No. 4,305,914. The process it describes attempts to obtain a jarosite precipitate with low non-ferrous metals content to make the jarosite more easily marketed.
Document WO 02/46481 A1 describes a procedure that appears to meet the requirements mentioned above, since it does not require a neutralizing agent for the iron to be precipitated as jarosite. This procedure follows the goethite process line, because in addition to neutral leaching it entails one or several stages of acid leaching followed by a reduction stage where the Fe+++ is reduced to Fe++ in the presence of the zinc concentrate, and a neutralizing stage during which calcine is used to neutralize partially or totally the acidity present in the solution. Finally, instead of continuing with the process that will result in the precipitating of iron as goethite, the jarosite is precipitated by means of injecting oxygen in the presence of sodium, potassium or ammonium ions, as well as a significant recirculation of jarosite solids under temperature conditions close to the boiling point of the solution. However, this procedure presents a series of difficulties that are probably the reason why it has not been possible to apply it to industrial processes. Indeed, it is clear that the working acidity during the jarosite stage depends on the amount of iron precipitated as jarosite, and therefore, the larger the iron concentration at the beginning of this stage, the more iron is precipitated, and subsequently, the greater the final acidity at which the iron is precipitated. In order to achieve an acceptable iron precipitation percentage working with high acidity levels it is necessary, on the one hand, to raise the working temperature to values close to the boiling point. This entails an unacceptable risk for people and facilities, unless autoclaves are used. On the other hand, as shown on the example provided in that same document, it is necessary to recycle the jarosite seed (although the document does not mention this must be done by recycling the underflow from the jarosite thickener) in significant amounts, which makes it necessary at this stage to also increase, significantly, the flow and the percentage of solids (according to the example shown in said document the circulating flow has to be increased during this stage by more than 100% with a high content of suspended solids). The consequences derived from the required operating conditions are high steam consumption, increasing the volume of the necessary equipment, and a considerable increase in the consumption of flocculant.
Proof of the lack of success achieved to date is that today none of these processes that attempted to improve the quality of the iron residue is being used and this residue continues to be stored in safety ponds, with the exception of a plant that generates hematite and those which use pyrometallurgical processes for treating the residues.
Patent Application PCT ES 2011/070265 describes a procedure similar to that described in document WO 02/46481 A1. The fundamental difference contributed by the process described in this document is that the maximum working acidity is limited, thereby limiting the admissible iron content of the initial solution and using non-polluting neutralizing agents according to the availability of those materials at the plant. This allows the jarosite precipitation stage to take place at lower acidities and temperatures without the need to recycle any jarosite seed. This procedure works well and is suitable for those plants that operate with zinc concentrates having low iron content (up to 5 of 6%), but it is not appropriate for treating zinc concentrates with high iron contents, since the procedure itself limits the maximum iron content in the solution that is treated during the iron oxidation and jarosite precipitation stage. This is relevant at this point in time, when the trend is to treat zinc concentrates with ever higher impurity rates, iron being the most abundant. Today, it is very usual to find zinc concentrates in the market containing between 8 to 12 percent of iron. In the daily practice this translates into the following: a plant working with zinc concentrates with an average 5% iron content generates a solution from the acid leaching stage containing between 18 to 20 g/l of iron. If they were to work with zinc concentrates containing an average of 9% iron content the iron content after the acid leaching stage would be 30 to 32 g/l. While in the first case the procedure described in document PCT ES 2011/070265 could be applied without any problem, in the second case applying that procedure would generate an excessively high acidity during the jarosite stage that would prevent the iron from precipitating efficiently, leaving a considerable amount of iron in the final solution. This would require a high amount of iron recirculating through all the leaching stages, as indicated later in the present document, unless the neutralizing agent would be used in quantities that are not usually available at production plants, which would also generate additional costs and the additional problem derived from using BZS as neutralizing agent, as it dilutes jarosite residue (therefore increasing its volume), a situation which is not recommended on economic grounds. It is for this reason that the procedure described limits the iron content in the solution resulting from the acid leaching stage to 25 g/l maximum. This condition, according to the procedure described, is only attained by limiting the iron content of the zinc concentrates treated, which entails an inconvenience for a good part of the zinc producers that would see their capacity for treating the concentrates available in the market limited by this condition.
As it is known in the industry, the plants working with the RLE system—the system more often used today—generate steam during the roasting process. This steam is later used to heat the solutions that are processed during some of the leaching and purification stages. The most important points of consumption for those plants that use the jarosite process are the acid leaching, jarosite precipitation and hot purification stages. It is important for the economic considerations of the process to manage judiciously the available steam because if more was needed the cost of generating would substantially increase the operational costs. Therefore, excessive flows and/or temperatures are to be avoided during the different leaching stages so the plant can be self-sufficient by using the steam generated during the roasting process.
It is well known in the art that precipitating iron as jarosite under atmospheric pressure conditions is an incomplete process because always a portion of the iron remains unprecipitated. In addition it should be taken into account the Fe++ present, which does not precipitate unless it is oxidized to Fe+++, In fact, on the one hand the oxidation of Fe++ to Fe+++, in acid medium is incomplete, so the non-oxidized portion of Fe++ remains in solution because jarosite is only formed from Fe+++, while on the other hand the Fe+++ present in the solution precipitates partially as jarosite depending on certain operational parameters such as acidity, temperature, residence time and the concentration of Na+, K+ or NH4+ ions used to form the jarosite. The presence of jarosite seed, recycling the underflow of the jarosite thickener may also affect the percentage of precipitated Fe+++ as claimed in patent document WO 02/46481 A1. The iron that does not precipitate as jarosite moves to the neutral leaching stage, returning again through all the stages of the leaching process until finally arriving to the iron oxidation and jarosite precipitation stage. The consequence of this behavior is a flow increase in all stages that will be larger the more the iron recirculates. This may affect the stability of the process as a whole, because first, during the acid leaching and jarosite precipitation stages steam is consumed to heat the solutions, and the larger the flow of these stages the larger the amount of steam needed. Currently there are some electrolytic zinc plants using the jarosite process that are encountering difficulties to maintain the stability of the leaching plant because they operate in such a manner that the iron recycled in the process is in the order of 50%. Therefore an objective of any process where iron is precipitated as jarosite must be that the precipitation be as complete as possible in order to minimize excessive iron recirculation.
Consequently, an objective of the present invention is to provide a hydrometallurgical method for recovering zinc in sulphuric media from sulphidic zinc concentrates having a high iron content that it will make possible to attain high rates of metal recovery.
Another objective of the present invention is to provide a hydrometallurgical method for recovering zinc in sulphuric media from sulphidic zinc concentrates having a high iron content in which an environmentally acceptable iron residue is obtained that can be used in other industrial processes, avoiding thus having to store it in safety ponds.
Another objective of the present invention is to provide a hydrometallurgical method for recovering zinc in sulphuric media from sulphidic zinc concentrates having a high iron content that manages efficiently the energy resources generated during the roasting process done at the electrolytic zinc plant to minimize operational costs.
Another objective of the present invention is to provide a hydrometallurgical method for recovering zinc in sulphuric media from sulphidic zinc concentrates having a high iron content that is capable of reducing iron recirculation through the various leaching stages to provide a stable and efficient operation at the electrolytic zinc plant.